Smelting of lead-containing ores



United States Patent SMELTING OF LEAD-NTAINING ORES Stephen WilliamKenneth Morgan and George Kenneth Williams, Bristol, England, assignors,by mesue assignments, to Metallurgical Processes Limited and TheNational Smeltiug Company Limited, doing business as MetallurgicalDevelopment Company, Nassau, Baharnas No Drawing. Application April 6,1954, Serial No. 421,454

Claims priority, application Great Britain April 10, 1953 18 Claims.(Cl. 7587) This invention relates to the smelting in a blast furnace ofmetalliferous materials containing lead and zinc and has for its objectthe provision of a blast-furnace smelting process for a wide variety ofmetalliferous materials containing lead and zinc values in which most ofthe lead and zinc are separately recovered in the form of metal.Throughout this specification and the appended claims metalliferousmaterials containing lead and zinc are to be understood as ores,concentrates, metallurgical products and any and all other metalliferousmaterials containing lead or zinc or both in amount susceptible of beingrecovered as metal in the blast furnace smelting process of theinvention.

To the best of our knowledge, no economical single pyrometallurgicalprocess for treating lead-zinc ores with the direct and separaterecovery of both lead and zinc as marketable lead metal and zinc metal,respectively, has heretofore been successfully practiced on a commercialscale. On the contrary, the successful pyrometallurgical treatment oflead-zinc ores has generally been possible only when the compounds ofthese two metals can be separated by some process of differentialconcentration, such as froth flotation. Even in favorable circumstances,such separation processes generally yield a zinc concentrate containingat least one percent of lead and a lead concentrate containing-severalpercent of zinc.

The pyrometallurgical method usually practiced for the production ofmetallic zinc from such zinc concentrates comprises distilling a mixtureof the oxidized concentrate with carbonaceous material to produce agaseous mixture of zinc vapour and carbon monoxide from which zinc iscondensed out as liquid metal. The heat required in this method isfurnished by electrothermic heating or by enclosing the charge inexternally heated retorts. In this method, lead is volatilized to anextent limited by the vapour pressure of lead at the temperature atwhich the gaseous mixture of zinc vapour and carbon monoxide leaves theheated charge. With vertical retorts, for example, the lead volatilizedis generally found to be about 0.1% to 0.2% of the weight of zincdistilled, being but little influenced by the amount of lead present inthe charge, provided that the charge contains suflicient lead to thussaturate the gaseous mixture. The remainder of the lead in the charge isleft in the retort residues and is lost when these are discarded asvalueless. Various processes, such as heating in rotating kilns, havebeen proposed and sometimes used for recovering the lead, together withany residual zinc, from zinc retort residues, but such processes areexpensive to operate and generally yield merely a mixture of zinc andlead oxides.

Lead concentrates are generally smelted in blast furnaces, which are sooperated that substantially all of the zinc goes into the slag as oxideor oxidized compound of zinc. By suitably controlling the fluxadditions, it is possible to operate lead blast furnaces so that theslag contains up to 25% zinc. The amount of zinc that can be toleratedin the charge to a lead blast furnace there fore depends on the weightof slag produced. With the amounts of gangue material usually present inthe ore and of fluxes customarily added, serious difficulties arise whenthe zinc content of the smelting charge is more than one quarter of thelead content. Zinc is thus a particularly undesirable impurity in thecharge to a lead blast furnace. The zinc present is lost when the slagis discarded as valueless. Sometimes such zinciferous blastfurnace slagsare treated by blowing pulverized fuel into the molten slag andrecovering zinc as zinc oxide from the hot gases thus generated, butthis process is expensive in fuel and depends very much for its economicfeasibility on finding a useful outlet for the large amount of heatcontained in the gases so generated. Such zinciferous slags have alsobeen treated by passing an electric current through the liquid slag,with a layer of coke on the surface thereof, and condensing metalliczinc from the gaseous mixture generated; to the best of our knowledge,this process becomes economically feasible only when the bulk of theslag to be treated is already available in molten form, so that inpractice the process is applicable only to the treatment of leadblast-furnace slags as they leave the furnace.

We have been actively associated in the development of an improvedprocess of smelting zinc ores in a blast furnace of which some of theoutstanding characteristic features are described in United StatesPatent 2,671,725 of Robson and Derham granted March 9, 1954, and thecopending application of Woods Ser. No. 374,322, filed August 14, 1953.Heretofore, no more lead has been tolerated in the zinc ore to beblast-furnace smelted by that process than heretofore customary in theaforementioned pyrometallurgical method for treating zinc concentrates.As the result of our investigations of the effect of increasing the leadcontent of the smelting charge in that process we have found, as indeedmight have been expected, that some of the lead reduced in smelting isvolatilized and some can be tapped from the bottom of the furnace. Inaddition, however, we have discovered that, in one respect, lead behavesmore favorably than might have been anticipated, in that the improvedzinc-smelting blastfurnace process takes lead in its stride, and nofurther addition of carbonaceous fuel is required for the reduction ofadditional lead oxide to lead metal, while the amounts of zinc andgangue materials are kept constant. On the other hand, We havediscovered that the presence of additional lead in the smelting chargecauses certain disturbances to the operation of the furnace and thecondenser, and one aim of the present invention is to overcome thesedisturbances by certain modifications of the operating conditions.

The most abundant ore of zinc is its sulphide, zinc blende. Azinc-smelting blast-furnace requires the zinc to be present mainly asoxide or oxidic compounds. Accordingly, sulphide ores are roasted beforebeing charged to the blast-furnace, this roasting preferably beingcombined with or followed by a sintering process which yields a productin a suitable physical condition for charging to the furnace. While suchroasting eliminates most of the sulphur, some sulphur is always stillpresent in the prodnot. In addition, carbonaceous fuels, such as coke,nearly -2,s1e,o22

is present in the slag tapped from the bottom of the furnace; whensufficient is present to saturate the slag, a sulphur-rich matte, ofwhich the main component is ferrous sulphide, separates from the slag.When copper and silver are present in the zinc ore, it is advantageousto have present sufficient sulphur to form a matte, since the copper andsilver are thus collected as sulphides in solution in the ferroussulphide.

When small amounts of lead compounds are added to the zinciferousblast-furnace charge, all the lead is volatilized and there is someincrease in the amount of sulphur volatilized. So long as the total leadintroduced is not greater than about one tenth of the weight of carbonburnt in the furnace, the additional amount of sulphur volatilized isnot large and it leads to no serious operational difficulties. Furtheradditions of lead, however, cause operational difficulties of two types.

Sometimes the introduction into the charge of lead amounting to 15% ormore of the weight of carbon burnt in the furnace causes no immediateoperational difficulties. Some lead metal is tapped from the bottom ofthe furnace. The amount of sulphur volatilized does not rise unduly.After some period of operation, generally a few weeks but sometimes onlya few days if lead and sulphur are present in the charge in very largeamounts, furnace pressures rise unduly. Opening up the furnace thenreveals that this is due to the formation of an accretion of interlockedacicular crystals of sulphide round the furnace offtake. This type ofbehaviour is often encountered with a relatively small pilot-scalefurnace using an invertedtrough oiftake, as described in the copendingUnited States patent application of Derham, Ser. No. 158,190, filedApril 26, 1950.

Sometimes the introduction of lead into the charge in amounts up to 25%or even more of the weight of carbon burnt leads to no lead being tappedfrom the bottom of the furnace. Instead, all the lead is volatilized andat the same time the amount of sulphur volatilized rises considerably. Astatistical study of a large number of data attained on a full scalefurnace shows that the ratio of additional sulphur volatilized toadditional lead volatilized, with this lead in the range between 10% andabout 30% of the weight of carbon burnt, corresponds quite closely tothe chemical equivalents of sulphur and lead, that is to say, in theratio of 32/207. This type of behaviour is encountered particularly witha full scale furnace working at a higher blasting rate per unit area andwith a rather coarser charge than is customarily used on the smallpilot-scale furnace. This associated volatilization of lead and sulphurcan sometimes proceed to the point at which nearly all the sulphur isvolatilized leaving in the extreme case only about 02-03% sulphur in theslag although the total sulphur present in the charge would besufficient to give 1.0-1.5% sulphur in the slag, all the remainder beingvolatilized.

In conjunction with our operations on a full-scale furnace andexperiments on a pilot-scale furnace, we have carried out laboratoryinvestigations to elucidate the mechanism of this transport of sulphurin the furnace. These show that when the charge to a Zinc-smeltingblastfurnace contains iron oxides as the predominant nonzinciferouscomponent and no free lime is present, the thermodynamically stablecondensed form in which any sulphur present can exist at temperaturesbelow about 1050 C. is zinc sulphide, whereas at higher temperaturesferrous sulphide becomes the stable form. If free lime is present thestable form of sulphur becomes calcium sulphide at temperaturesprevailing in any zone of the blastfurnace. The temperature at which thegases leave the furnace charge is generally between 950 C. and 1000 C.,typically 975 C. At 975 C. the vapour pressure of lead is such that in asaturated lead vapour the weightof lead is equal to about 7% (thecorresponding figures being approximately for 950 C. and for 1000" C.)of the weight of carbon contained in the carbon monoxide sequentlyrecovered in the water scrubbers.

4 and carbon dioxide, this weight of carbon corresponding,

of course, to the Weight of carbon burnt in the furnace. In the reactionZnS-f-Pb (liquid)=Zn (gas) +PbS (gas) the equilibrium at 950-1000 C. issuch that with zinc present in a volume of concentration of 5-6%, thevolume concentration of lead sulphide in the gas will be about /5 of thevolume concentration of lead vapour; that is to say, at equilibrium somelead sulphide. but in relatively small amounts, would leave the furnaceas gas. At the bottom of the furnace where the temperature may be 1100C. or higher, the equilibrium between liquid lead and ferrous sulphidewith metallic iron is such that a considerable concentration of leadsulphide can be present at equilibrium; if free lime is presentequilibrium is attained here with the formation of calcium sulphide andthe presence in the gas of a lower but still appreciable concentrationof lead sulphide, and the lime in the upper parts of the furnace canreact with this lead-sulphide vapour so that the amount of sulphurfinally volatilized from the furnace becomes greatly reduced.

In the light of equilibrium studies it can be seen if the amount of leadadded to a charge does not exceed about 7% of the weight of carbonburnt, all this lead can be volatilized as lead vapour and will tend tocarry with it only a small concentration of lead sulphide. Once a largeramount of lead is added to the furnace not all of it can be volatilizedfrom the upper zones of the furnace. Some liquid lead goes down to thehotter zones of the furnace. Here it reacts with zinc sulphide, ironsulphide or any other sulphur compounds present to form a considerableconcentration of lead sulphide vapour. Passing up to the furnace thislead sulphide vapour, if it attains equilibrium with the zinc vapour inthe upper part of the furnace, mostly reacts according to the equationPbS (gas)+Zn (gas)=Pb (liquid) +ZnS (solid) Some of this zinc sulphidemay be deposited as an accretion. On the other hand, under certainoperating conditions this lead sulphide vapour may escape reacting withzinc vapour in the furnace and be transported as such to the condenser.Here it reacts to give presumably very small droplets containing zincsulphide and lead in intimate admixture. The mixed droplets of zincsulphide and lead are not readily carried down by the spray of leadthrown up in the condenser; and act as nuclei for the deposition of moreliquid zinc, which in turn cannot readily be trapped in the condenser.As a consequence more dross is formed in the condenser. In so far as thedroplets are carried forward they are recovered, with their associatedcondensed zinc, in the blue powder subindced, a statistical study of anumber of operating periods on a fullscale furnace has shown that eachadditional 1.0 lb. of sulphur volatilized from the furnace causes theformation of an additional 14.8 lbs. of dross and blue powder,

containing 5.5 lbs. of lead and 6.1 lbs. of zinc.

In the light of these facts we have found it necessary when operatingwith mixed lead zinc ores to take special precautions to eliminate thesulphur more completely during the roasting than would have beenotherwise regarded as necessary. A statistical study of the resultsshows that the harmful effect of sulphur is proportional to the amountpresent. Within the ranges that can practically be attained there is nolower limit below which the sulphur present can be regarded as not beingharmful. So long. however, as the sulphur is below 0.8%, the effect isnot unduly serious. We regard the 1.5% of sulphur in the metalliferousmaterials charged as the upper limit to be tolerated to obtainreasonable operating conditions.

It has been customary to add some lime to the blast furnace chargetoobtain a slag of suitable composition.

From the point of view of slag composition it does not matter in whatform the lime is introduced and experiments were made with adding thislime as a slag, the lime being already combined as calcium silicate. Itwas discovered, however, that the amount of sulphur volatilized wasreduced by having free lime as such present in the charge. The reasonwhy this was so was explained by the aforementioned laboratoryinvestigations, that is, the formation of calcium sulphide. Whensmelting mixed lead-zinc ores therefore, it is preferable to have freelime present in the charge. It has been found that even if the lime isincorporated with the ore before it is sintered or sinter-roasted, muchof the lime is still present in the free state or at least in the formof some compound which is substantially as active as free lime withrespect to reducing the volatilization of sulphur from the furnace.

Another procedure that was found helpful in reducing the amount ofsulphur volatilized was to introduce some metallic iron in the charge.This was more effective in helping to trap sulphur than the oxidizediron compounds which are nearly always present. In general, however,this is a less preferred method than the use of lime, both because thelime is somewhat more effective and because addition of lime is oftenrequired as a slag component, Whereas to form a suitable slag it is notgenerally desired to add more iron.

With mixed lead-zinc ores it is undesirable to have sufficient sulphurreaching the bottom of the furnace to form a sulphide matte. Analternative procedure must therefore be followed to recover any silverand copper present. We find that when lead metal is tapped from thebottom of the furnace, this dissolves nearly all the silver and most ofthe copper, so that these metals can still be recovered.

It was thus established that, contrary to expectations, it was importantto keep the sulphur content of the smelting charge as low as possible,since the joint presence of lead and sulphur produced unanticipatedobjectionable effects in both furnace and condenser.

It was further found that the lower the temperature at which the gasesare withdrawn from the smelting charge, the lower is the amount of leadsulphide vapour carried by the gases, and therefore the lower the amountof zinc sulphide fume present in the condenser. The gases at the pointat which they last make contact with the furnace charge must be at atemperature at which the zinc vapour contained therein cannot react withcarbon dioxide to produce Zinc oxide according to the reaction,

The equilibrium constant in this reaction is such that, for instance, agas containing by volume 6% zinc vapour, 7% carbon dioxide and 25%carbon monoxide can have a continued stable existence at temperaturesonly above 965 C., while a gas containing 6% zinc vapour, 10% carbondioxide and 22% carbon monoxide can have a continued stable existenceonly above 1000 C. In order to reduce the amount of lead-sulphide vapourcarried from the furnace, it is desirable that the gases should leavethe furnace charge at a temperature only slightly above the equilibriumtemperature at which Zinc Vapour could react with carbon dioxide toproduce zinc oxide.

. It was also found that in the previously preferred form of invertedtrough gas offtake, below the top level of the smelting charge, thegases were superheated slightly above their. equilibrium temperature. Bytaking the gases off from the top of the charge, it was found that thelead sulphide content of the gases was reduced, since at this point thegases are cooled to approximately the equilibrium temperature for thereversible reaction But while this alleviates the sulphur trouble itaccentu- 6 ates the tendency for zinc vapour to oxidize during thetransfer of the gases to the condenser, as hereinafter pointed out.

In operating a zinc blast furnace the ratio of Zinc volatilized to thecarbonaceous material consumed depends on the heat balance, which can,of course, be improved if operating conditions are so controlled thatmore air is consumed per unit of carbon and hence a greater proportionof the carbon is oxidized to carbon dioxide instead of to carbonmonoxide. We have found, however, that it is impossible to generate agas containing a very large amount of carbon dioxide and at the sametime to attain a good zinc elimination. We find that when good operatingconditions are being attained the volume percentage of carbon dioxide inthe gas leaving the charge does not greatly exceed that of the zincvapour; typically the gas leaving the charge contains about 5% zinc andabout 6% carbon dioxide, and after admixture of air or oxygen thecontent of carbon dioxide may rise to about 7%.

The reason why the gas leaving the charge should contain zinc vapour andcarbon dioxide in a volume concentration not greatly exceeding that ofthe Zinc vapour can be explained most readily in relation to a furnacein which the gases are withdrawn from the top of the charge and thewhole of the air blast is introduced at the bottom of the furnace, sothat flow of charge and gases is countercurrent throughout the furnace.

Analogously with other metallurgical blast furnaces the gas generated atthe bottom of a zinc-smelting blast furnace contains carbon monoxide,which in its passage up the furnace can reduce zinc oxide according tothe equation ZnO+CO=Zn+CO This reaction absorbs a large amount of heat,so that, if carbon monoxide constituting 1% by volume of the gas thusreacts with Zinc oxide to give 1% by volume of zinc vapour in the gas,the heat required with countercurrent flow of charge and gas will reducethe temperature of the gas by about 50 C. The temperature at which thecharge components begin to melt may be in the region of 1100-1150 C.;the gases finally leave the furnace charge at not much below 1000 C.Consequently only about 3% by volume of zinc vapour can be formed byreduction of Zinc oxide by carbon monoxide below slag-meltingtemperatures. Once a slag has been formed it takes zinc oxide intosolution, and such zinc oxide dissolved in slag is more difficult toreduce than free zinc oxide; also, some of the zinc may initially beintroduced as difiicultly reducible compounds, such as silicate oraluminate; to ensure that as much as possible of such zinc is reducedand liberated as zinc vapour, it is necessary that the gaseousatmosphere there should be as reducing as possible; therefore conditionsmust be so controlled that the air blast introduced at the bottom of thefurnace reacts with the carbonaceous material to give initially mostlycarbon monoxide with relatively little carbon dioxide. As the gasestravel up the furnace the carbon monoxide reacts with zinc oxide andother readily reducible oxidic Zinc compounds, according to theequation,

thus producing equal volumes of zinc vapour and carbon dioxide. Duringthe travel of the gases up the furnace, some carbon dioxide reacts withcarbon according to the equation,

but this reaction takes place to only a small extent. Consequently, theoverall carbon dioxide content does not greatly exceed that of the zincvapour.

We have discovered that the zinc vapour in the blastfurnace gases can beefficiently condensed to zinc metal with a much higher content of carbondioxide in the gases than heretofore thought possible. This opened up 7new possibilities in the blast-furnacesmelting of lead-zinc ores with aneconomical consumption of fuel which we have further discovered, as aresult of our investigations. Simplified, the operation of the lead-zincblast furnace may be regarded as depending on the generation of carbonmonoxide in the furnace and these two reactions:

The first reaction is exothermic and therefore is a source of heat. Thesecond reaction is highly endothermic. We have found that a primaryprerequisite for commercially successful lead-zinc blast-furnacesmelting is maintenance of a reducing environment throughout thesmelting charge at a temperature sufliciently high to reduce andvolatilize zinc with an amount of carbonaceous material in the chargenot substantially greater than that needed to provide the heat requiredfor smelting the zinc. Good zinc elimination is thus attained, and thecarbon dioxide content of the blast furnace gases may be increased (bythe reduction of lead oxide mostly above and partly in the slag) toaround 11% without objectionably impairing the condensation of zincvapour to zinc metal.

Based on the foregoing discovery and observations the present inventioncontemplates improvements in the method of operating a blast furnace inwhich a preheated charge containing oxidic zinc materials andcarbonaceous fuel is introduced into the top of the furnace, preheatedair is blown into the bottom of the furnace, molten slag is tapped fromthe bottom of the furnace and a gas containin zinc vapour, carbonmonoxide and carbon dioxide is withdrawn from a higher part of thefurnace, in accordance with which a substantial amount of oxidic leadcompounds is incorporated in the charge and the amount of combustiblecarbon charged is regulated so that the heat generated by combustion ofthis carbon in the preheated air to produce a mixture of carbon monoxideand carbon dioxide in such a ratio that this gaseous mixture is highlyreducing to the oxidic zinc and lead compounds present in the slaginitially formed in the furnace, together with the heat generated by theexothermic reaction between the oxidic lead compounds and carbonmonoxide to produce lead metal and carbon dioxide, is substantially nomore than that necessary (a) for compensating for heat losses from thefurnace, (b) for reducing the oxidic zinc compounds endothermically bycarbon monoxide to produce zinc vapour and carbon dioxide, (c) formelting substantially all of the charge components other than zincoxide, lead oxide and carbon, and maintaining the resultant slag at suchan elevated temperature that substantially all the zinc and lead oxidesdissolved therein are reduced to liberate zinc vapour and lead metal,(d) for raising the temperature of the gases in the zone of the furnacewhere they last make contact with the furnace charge to a point at whichthe zinc vapour contained therein cannot react with the carbon dioxidecontained therein to form undesired zinc oxide, and (e) for volatilizingsome of the lead metal formed, the amount of lead volatilized not beingsubstantially above that unavoidable minimum amount required to saturatethe gases leaving the charge with lead vapour at the temperature atwhich these gases finally leave the charge; and withdrawing molten leadfrom the bottom of the furnace.

The smelting charge and the air blown into the furnace are preferablypreheated to temperatures of at least 600 C. and 500 C., respectively.Alternatively, the carbonaceous material (e. g. coke) of the smeltingcharge may be separately preheated, in which case the carbonaceousmaterial may be more highly heated (say 800l000 C.) while the othercharge materials (preferably sintered) are heated to a lower temperature(say SOD-750 C.), the separate heating temperatures being correlated toimpart to the charge as a whole an average or mean temperature between600 and 900 C. when introduced into the blast furnace. Suchseparatepreheating-of thecar- "8 bonaceous material and other chargematerials is advantageous where the latter are sintered and the leadcontent is relatively high, since lead lowers the temperature at whichthe sinter tends to soften.

As hereinbefore mentioned, the reduction of zinc oxide by carbonmonoxide is a highly endothermic reaction, but the reduction of leadoxide by carbon monoxide is an exothermic reaction and the heat evolvedin the latter reaction beneficially contributes to the maintenance ofthe necessary highly heated smelting charge. As a consequence ofpreheating the charge and the air blast, a relatively large amount ofextraneous heat is introduced into the furnace. The preheated air blastin conjunction with locally consumed coke provides the necessary largeamount of heat locally required to maintain a relatively hightemperature in the furnace hearth. The smelting charge should containsufficient zinc to justify a condensing stage for zinc vapour in theblast furance gases and should contain lead in such amount that moltenlead metal is tapped from the bottom of the furnace along with a slag oflow zinc content. As already explained, no fuel is required for thereduction of the lead oxide. The amount of lead that can be reduced isgoverned by the consideration that the total carbon-dioxide content ofthe gas should not greatly exceed 11% by volume. This limit is generallyattained when the weight of lead introduced is about 2.5 times theweight of carbon burnt.

As hereinbefore explained, the fuel value of the carbon burnt, inconjunction with the sensible heat introduced with the air blast andwith the solid materials charged, is used for a number of purposes, themost important of these being for counteracting heat losses from thefurnace, for melting as a slag the materials other than lead oxide andzinc oxide present in the charge (including the coke ash), for reducingthe zinc oxide and for raising the temperature of the gaseous productsof the reaction; compared with the foregoing items, the amount of heatrequired for volatilizing the relatively small amount of lead needed tosaturate the gases leaving the furnace charge with lead vapour iscomparatively small,

and, in general, the amount of heat required to reduce and volatilizethe quantities of other volatile metals, such as cadmium, that may bepresent is negligibly small. The ratio of fuel burnt to zinc volatilizedtherefore depends on a number of factors, such as the temperatures towhich the air blast and the charge are preheated, the heat loss from thefurnace and the amount of slag forming materials present. On a typicalfurnace, with charge preheated to 800 C. and air blast to 600 C., thecarbon consumption may be calculated as the sum of of the weight of zincto be volatilized, and 20% of the Weight of slag to be formed. With ahigh-grade mixed zinc-lead ore in which the weight of slag formed may be70% of the weight of zinc present, this means that the carbon consumedmight be about 104% of the weight of zinc reduced and volatilized. Forreasons hereinhefore discussed, the sulphur content of the smeltingcharge should not exceed 1.5%, and preferably should not exceed 0.8%.

The gaseous product resulting from the smelting operation (blast furnacegases) is withdrawn from the furnace and transferred through a suitableflue system to a shock-chilling condenser, such as a lead-splashcondenser, where about 90% of the zinc vapour in the gases is condensedand recovered as molten zinc metal. The blast furnace gases contain arelatively large volume of nitrogen (e. g. 61-63%), a relatively smallervolume of carbon monoxide (e. g. 24-27%), a relatively small volume ofzinc and lead vapours (e. g. 56%), and a volume of carbon dioxidecorresponding approximately to that generated by reduction by carbonmonoxide of the zinc oxide and lead oxide present in the charge (e. g.610%). In practice, the volume of lead vapour in the blast furnace gasesWill be around 3% of the volume ,of zinc vapour therein, and the balanceof the lead in cluded in the smelting charge, exclusive of the verysmall amount of lead sulphide that accompanies the furnace gases, istapped from the bottom of the furnace as molten lead metal along withmolten slag or" low zinc content. By weight, the amount of lead vapourin the blast-furnace gases is about one-tenth that of the zinc vapour.With lead present in the smelting charge in amount chemically equivalentto the zinc (i. e. 207.2/65.4=3.15 times as much lead as zinc byweight), the blast-furnace gases will contain, by volume, about zincvapour and about carbon dioxide, with about 0.16% lead vapour, whichvolume of lead vapour represents about 3% of the total lead in thesmelting charge, the remainder of the lead in the charge being recoveredas molten lead metal from the bottom of the furnace.

Not only must the blast-furnace gases be highly reducing in character,to prevent oxidation of the zinc and lead vapours, but the gases must bewithdrawn from the blast furnace at a sufficiently high temperature topermit their travel to the shock-chilling zone of the condenser withoutappreciable oxidation of zinc vapour.

As hereinbefore explained, to reduce the amount of lead sulphide carriedin the furnace gases it is desirable that the gases should leave thefurnace charge at such a temperature that a further small fall intemperature will permit the zinc vapour and the carbon dioxide containedin the gases to react to form Zinc oxide, and in practice we have foundit advantageous to introduce a controlled amount of an oxygen-containinggas, such as air, oxygenenriched air, or oxygen, into the blast furnacegases being withdrawn from the furnace and thereby heating the gases bythe resulting oxidation of carbon monoxide therein to a temperaturesubstantially above their initial normal temperature and sufficientlyhigh to permit their travel to the shock-chilling Zone of the condenserWithout appreciable oxidation of zinc vapour, as more fully described inthe aforementioned Woods application Serial No. 374,322.

The invention will now be described, by way of example, as applied tothe treatment of a mixed lead-zinc sulphide ore containing at least 5%of lead, where all of the smelting charge ingredients, other than thecarbonaceous material, are sintered with simultaneous roasting of thesulphite ore.

The blast-furnace charge is made up of coarse porous sinter andcarbonaceous material (usually coke and for convenience so hereinreferred to). The sinter contains all of the smelting charge ingredientsother than the coke. In addition to blending all of these ingredients,sintering also serves to roast the raw or fresh lead-zinc sulphide oreand to calcine such limestone as is required to furnish the lime whichmay serve the dual purposes of preventing undue volatilization ofsulphur from the blast furnace and of giving a slag of the desiredcomposition. The fuel required for sintering is the combustible sulphurin the lead-zinc sulphide ore, and since the optimum fuel requirementfor sintering is a determined factor, the amount of raw ore iscorrelated with respect to the sintering charge as a whole. Thus, theusual sintering charge preferably contains from 6% to 7% by Weight ofsulphur, but the sulphur content may vary from about 4% to as high as9%, depending in some measure upon the character of the other chargeingredients; in particular, the higher the lead content of the charge,the lower is the required sulphur content of the sintering charge. Thesulphur content of the ore is usually between 14% and 33%, the lowersulphur contents generally occurring with ores containing the higherratios of lead to zinc. In making up a sintering charge, the amount ofother ingredients mixed with the ore is usually from 2.5 to 5 times asgreat as the amount of raw ore. The bulk of such other chargeingredients may be fines from a previously sintered similar charge, leadblast-furnace slags of high zinc oxide content (c. g. 10 to and othersuitable materials contain- 10 ing zinc and/or lead in amount warrantingblast-furnace smelting.

Other sintering charge ingredients are fluxing agents such as limestoneand sand (silica) in amounts to give a slag of suitable composition, anddross and blue powder returned from the stage of zinc vapourcondensation. To obtain good zinc elimination in lead-zinc blastfurnacesmelting, we have formed a slag of relatively high melting point, andfavorable proportions of ferrous oxide (FeO), lime (CaO) and silica (SiOin such a slag are, respectively, about 1:l.5:1.5. While of high meltingpoint, such a slag is fluid and free-flowing in our practice Where thetuyeres of the blast furnace are supplied With preheated air. Theinclusion of the fiuxing agents in the sintering charge assures moreintimate mixing thereof with the other ingredients of the charge withattendant improvement in the smelting reactions in the blast furnace.The inclusion of limestone in the sintering charge is advantageous andimportant, since if added as such to the blast-furnace charge, itscalcination in the blast furnace consumes so much heat near the top ofthe smelting charge so as to make it impracticable to attain andmaintain the contemplated high temperature of the blast-furnace gasesabove the top level of the charge. Furthermore, the inclusion oflimestone as such in the smelting charge is undesirable because of thelarge amount of carbon dioxide which its calcination would introduceinto the blast-furnace gases. As noted above, some of the lime should bepresent as free lime, or in some other active form, in order to minimizethe volatilization of lead sulphide.

Roughly, about of the zinc vapour entering the lead-splash condenser iscondensed and recovered as molten zinc metal. The remaining 10% of thezinc vapour is recovered in the dross periodically removed from thecondenser and in the blue powder washed out of the condenser exhaustgases. Both the dross and blue powder contain lead, some of which isentrained from splashing molten lead in the condenser, while some comesfrom lead vapour in the blast furnace gases depending to some extent atleast upon the relative amounts of lead and zinc in the sinter chargedinto the blast furnace. In our present usual lead-zinc blast furnacesmelting practice, the combined dross and blue powder customarilycontains by analysis about 35% Zinc and 35% lead. Such lead vapour asmay be condensed in the lead-splash condenser is absorbed by thecirculating molten leadzinc metal of the condensing system, and itsaccumulation therein may require bleeding-off of some of the moltenlead-zinc metal in the system from time to time. The amount of returneddross and blue powder included in the sintering charge is approximately4% by weight based on the dry weight of the sintering charge.

The various sintering charge ingredients are delivered in predeterminedamounts to a mixer where sufiicient water (not less than 5% and usuallyaround 6% on the dry Weight of the other ingredients) is added to form amass that can be satisfactorily handled in a mixer. After thoroughmixing, the sintering charge is fed to a sinter machine, such as aconventional down-draft sinter machine, and the exhaust gas of themachine is delivered to a sulphuric acid plant.

The sinter made on the machine should be porous and low in sulphurcontent. Sulphur in the sinter should not exceed 1.5% and preferably isless than 0.8 by Weight, for the reasons hereinbefore discussed. Thesinter is discharged from the machine onto a screen, conveniently of thebar or grizzly type, where the coarse sinter is separated from the finesinter. Where insuflicient sinter fines are obtained in screening tomeet the amount required in making up the sinter charge, some of thecoarse sinter may be crushed to provide the required amount of finesinter. The coarse sinter is of suitable size for blast furnacepractice, and to this end most of it should be over l inch but less than3 inches in size.

Coarse sinter and coke in suitable proportions'for smelting aredelivered to a preheater. In general, with coke initially containingaround 8% water and having an ash content of about 8% (on the drybasis), the coke consumption (on the initial wet basis) is about 1.1 tonper ton of zinc present plus about 25% of the weight of the othersubstances in the sinter other than zinc oxide and lead oxide, and inthe present usual practice is approximately of the order of 40 parts ofcoke for 100 parts of coarse sinter containing around 30% zinc and 30%lead. The preheated charge, usually at a temperature around 850 C., isintroduced into the top of the blast furnace through a suitably sealedcharging device. The air introduced through tuyeres near the bottom ofthe blast furnace is preheated, say to a temperature around 600 C. andpreferably as high as practicable. Molten slag and lead are tapped fromthe bottom of the furnace. The blast furnace gases accumulate in theupper part of the furnace above the level of the charge at a normalinitial temperature of around 970 C., and of an initial compositionusually about as follows:

Percent Zinc 5-6 Lead 0.1-0.2

Carb on dioxide 6-8 Carbon monoxide 2427 Nitrogen 61-63 A controlledamount of an oxygen-containing gas, such as air, oxygen-enriched air, oroxygen, is introduced into the blast furnace gases accumulating abovethe level of the charge, in order to raise the temperature of the gases,say to at least 1000 C., by the oxidation of carbon monoxide, as morefully explained in the aforementioned Woods application Ser. No.374,322. After such heating, the blast-furnace gases will contain about1% more carbon dioxide and about 2% less carbon monoxide, while the zincand lead concentrations may be slightly reduced owing to dilution by anynitrogen that has been introduced. Air, because of its high nitrogencontent, should be preheated to a temperature around 600 C., andpreferably as high as practicable, when used as the oxygencontaininggas, but preheating of oxygen (used as the oxygen-containing gas) isunnecessary, since it contains no diluent and its cooling effect uponthe blast furnace gases is negligible. The oxidation of carbon monoxideby the oxygen content of the oxygen-containing gas sufficiently raisesthe temperature of the blast-furnace gases (e. g. to at least 1000 C.)to permit their travel and delivery to the shock-chilling zone of thecondenser at a temperature above that at which any appreciable amount ofzinc vapour reacts with carbon dioxide in the gases.

The blast-furnace gases, heated as hereinbefore described to atemperature substantially above their normal initial temperature, aretransferred through a suitable flue system to a condenser of theshock-chilling type, such for example as the two-stage lead-splashcondenser and associated molten zinc metal separator described in theaforementioned U. S. Patent 2,671,725 of Robson and Derham. The fluesystem is of the heat-insulated type to minimize any substantial loss ofheat in the gases in the course of their travel from the blast furnaceto the shockchilling zone of the condenser, so that, even though thecarbon dioxide content of the gases may have been raised to around 10%by the aforementioned oxidation of carbon monoxide, there is practicallyno tendency for zinc vapour to react with carbon dioxide in the gases.

Due to the relatively high concentration of carbon dioxide in theblast-furnace gases, zinc vapour will react with carbon dioxide in thegases to form zinc oxide at temperatures below about 950 C. Therefore,in order to recover the zinc in metallic; form the gases should be 12practically instantaneously cooled through the temperature range of fromabout 950 C. to about 650 C. This is effected for all practicalpurposes, in the aforementioned lead-splash condenser, by bringing thehot gases, at a temperature not lower than 950 C., into intimate contactwith a shower of molten lead whose temperature is lower than 650 C.(usually around 570 C.), whereupon zinc vapour condenses and dissolvesin the molten lead. Thus, immediately upon their delivery to thecondenser, the blast-furnace gases are brought into shockchillingcontact with molten lead in the first stage of the condenser. Herecondensed zinc vapour dissolves in the molten lead and the Zinc-richmolten lead is withdrawn from the condenser at a temperature below 650C. (usually around 570 C.) by a pump to cooling equipment,

such for example as the water-cooled trough system described in UnitedStates patent application of Keeping Ser. No. 361,041, filed June 11,1953. The cooled molten lead, from which molten Zinc separates as theleads saturation point for zinc decreases in the course of cooling, isconducted to a separator, where at a temperature of around 450 C. themolten zinc floating on the zinc-depleted molten lead is overflowed to acollecting ladle or the like. The zinc-depleted molten lead at atemperature of about 450 C. is returned to the second stage of thelead-splash condenser, where the molten lead picks up residual zinc inthe gases coming from the first stage of the condenser, with anattendant rise in temperature of the molten lead in this stage to about470 C., and at about the latter temperature molten lead is conducted tothe first stage of the condenser. In practice, the Zinc-rich molten leadwithdrawn from the first stage of the condenser at a temperature ofabout 570 C. will contain about 2.5% zinc, and the zinc-depleted moltenlead returned to the second stage of the condenser at a temperature ofabout 450 C. will contain about 2.2% zinc. In both stages of thecondenser, the flow of the gas and molten metal is continuous andcounter-current.

The resdiual zinc in the gas exhausted from the second stage of thecondenser is removed, mostly in the form of blue powder, in waterscrubbers. This blue powder contains a substantial amount of lead, someof which arises rom minute drops of molten lead entrained by the gasstream as it passes through the splashing molten lead in the condenser,and some arises from minute droplets which have been formed bycondensation from lead vapour leaving the furnace and have failed to betrapped by the splashed molten lead in the condenser. After settling,the wet blue powder is dried and returned to the sintering charge.Dross, mostly in the form of zinc and lead oxides, is formed andaccumulates in the condenser, and is periodically removed, usually bytemporarily shutting down the condenser. In practice, the amount ofdross slightly exceeds the amount of blue powder.

By a slag of low zinc content we mean that the slag generally containsnot more than about 5% of the zinc content of the initial smeltingcharge introduced into the blast furnace, which represents about 95%recovery of zinc as vapour in the blast furnace gases. However, therecovery of Zinc as vapour in the blast furnace gases is influenced bythe ratio of slag-forming materials (that is, substantially all of thecomponents other than zinc oxide, lead oxide and carbon) to zinc in thesmelting charge, and, for example, when a sinter-roasted mixture of leadconcentrates containing 75% lead with a slag containing 17% zinc isbeing treated the recovery of zinc as vapour may be only between andthis recovery being economically satisfactory when the zinciferousmaterial charged is a slag of low Zinc content. The actual zinc contentof the slag, by analysis, will depend upon the amount of slag-formingmaterials present in the smelting charge and coke and may vary from aslow as 1% up to 10%.

The invention provides an, economical processfor sepl l l l aratelyrecovering zinc metal and lead metal from zinc ores of high lead contentas well as from other metalliferous materials containing lead and zincby a single blast furnace smelting operation. The blast furnace chargecontains sufiicient zinc to warrant the recovery from the blast-furnacegases of zinc metal. The lead content of the blast-furnace charge is notcritical, but is at least suiiicient to warrant recovery of molten leadmetal from the molten product of the furnace and generally is higherthan the zinc content and may in some cases be slightly more than threetimes the Zinc content, by weight. Conveniently, the molten product maybe withdrawn from the bottom of the furnace into a pot or ladle where,after settling, the supernatant slag and settled lead metal may beseparated in any suitable manner. Alternatively, the blast furnace maybe provided with a fore-hearth for collecting molten lead metal. Themolten lead metal may contain silver, gold, copper, tin, antimony andbismuth, where the smelting charge contains any of these metals inappreciable amounts. The molten zinc metal recovered from thecondensation stage will contain lead, between 0.9% and about 2%, and maycontain cadmium and arsenic and part of the tin, antimony, and bismuthwhere present in appreciable amounts in the smelting charge.

Because of the aforementioned relative amounts of zinc and lead vapoursin the blast furnace gases, the smelting charge (excluding coke) mustcontain lead and zinc in a ratio larger than 1:10 if any molten leadmetal is to be formed at the bottom of the furnace. Since, in ourlead-zinc blast furnace process, the smelting charge (excluding coke)seldom contains more than 50% Zinc, it must contain more than 5% lead,and for practical economic operation should contain at least lead. Thelimits to the ratio of lead to zinc in the smelting charge dependsomewhat on the amount of slag-forming material present in the charge,and can be illustrated by the following two examples:

With a smelting charge obtained by sintering a highgrade zinc blendewith relatively small amounts of flux additions, it has been found thata blast-furnace gas can be generated containing, by volume, about 6%zinc, 7% carbon dioxide, 25% carbon monoxide and 62% nitrogen. Thecomposition of this gas, it may be noted, implies the consumption of5.33 atoms of carbon per atom of zinc, which is equivalent to about 0.98lb. carbon per lb. of zinc. Above the charge level in the furnace, someair (3% of the gas volume) is introduced into the gas, and theapproximate composition becomes, by volume, 6% zinc, 8% carbon dioxide,23% carbon monoxide and 63% nitrogen.

If now lead oxide is added to such a charge, the ratio of carbon to zincbeing kept the same, some of the carbon monoxide will be used inreducing the lead oxide. If the weight of lead so reduced is 1.6 timesthe weight of zinc volatilized (the atom ratio of lead to zinc being0.5), the volume of carbon monoxide used in reducing the lead will be 3%of the total gas volume, so that the compo sition of the blast-furnacegas as it leaves the charge will be, by volume, 6% zinc, 10% carbondioxide, 22% carbon monoxide and 62% nitrogen. After introducing air (3%of the gas volume) into the blast-furnace gas, the volume composition ofthe gas will become 6% zinc, 11% carbon dioxide, 20% carbon monoxide and63% nitrogen.

Where the process is practiced for sinter-roasting and blast-furnacesmelting of solely a lead-zinc sulphide ore, the raw sulphide oreincluded in the sintering charge will be the lead-zinc sulphide orecontaining at least 5% of lead, and the bulk of the other materialincluded in the sintering operation will be sinter fines containing morethan 5% lead. Due to the inclusion of dross and blue powder in thesintering charge, the sinter will always contain substantially more than5% lead when the lead content of the fresh lead-zinc ore is about 5%,and when 14 treating the lead-zinc ores of higher lead content thesinter will contain correspondingly higher amounts of lead.

Lead blast furnace slags can be advantageously smelted in accordancewith the invention. Such slags typically contain about 1418% zinc. Toattain good elimination of Zinc from such slags, such a ratio of fuel toslag is employed that the blast-furnace gas contains only about 3% zincvapour by volume, with 3% carbon dioxide, 31% carbon monoxide and 63%nitrogen. After introducing 3% by volume of air into the blast-furnacegas above the charge level, the volume gas composition becomes about 3%zinc, 4% carbon dioxide, 29% carbon monoxide, 64% nitrogen. It isaccordingly feasible to add to the smelting charge lead oxide so that 2atoms of lead are present per atom of Zinc, whereby the weight of leadpresent in the charge is about 6.3 times the weight of zinc. Thecomposition of the blast-furnace gas, by volume, is 3% zinc, 9% carbondioxide, 25% carbon monoxide, and 63% nitrogen. After introducing 3% byvolume of air into the blast-furnace gas its composition becomes 3%zinc, 10% carbon dioxide, 23% carbon monoxide, and 64% nitrogen. Fromsuch a gas the zinc can be condensed in a lead-splash condenser, whichindeed can deal with gases containing down to 2% zinc by volume. With 3%of Zinc vapour in the blast-furnace gas, the carbon consumption is 11 /3atoms per atom of zinc, or about 2.1 lbs. of carbon per lb. of zinc.With 2% of zinc vapour in the blast furnace gas, the carbon consumptionwould be about 3.1 lbs. per lb. of zinc. It is because the ratio ofcarbon consumption to zinc becomes uneconomically high as the zincvapour concentration in the blast-furnace gas decreases that treating agas containing less than 2% zinc is not warranted, rather than becausethere is any definite failure of the lead-splash condenser at such lowzinc concentrations.

The invention permits the inclusion of lead oxide in a zinc-smeltingblast-furnace charge until the reduction of lead oxide by carbonmonoxide causes the carbon-dioxide content of the blast-furnace gas torise to any point not exceeding about 11% by volume. The permissibleinclusion of lead oxide, in relation to the zinc in the charge, istherefore particularly high when the Zinciferous material is a low-gradeproduct such as a blast-furnace slag, where, in the absence of leadoxide, the zinc content of the gas would be low and the'carbon-dioxidecontent would also be low.

A smelting plant that has been separately smelting zinc and leadconcentrates obtained by froth flotation from a mixed sulphide ore willgenerally have accumulated a stock of a zinc-containing leadblast-furnace slag. Such a plant, adopting the present invention, maynow directly treat the mixed sulphide ore, by sinter-roasting, withaddition of an appropriate amount of lead blast-furnace slag to thesmelting charge, thus gradually working off the dumps of this material.The slag may be included in the sintering charge or separately preheatedand added to the blast furnace.

A smelting plant that has been treating only lead concentrates in ablast furnace will also have accumulated the same type of zinciferousblast furnace slag. By the present invention, such a plant may recoverzinc from the zinciferous slag and simultaneously smelt fresh leadconcentrate. Although the lead blast-furnace slag will contain a littlelead and the lead ore a little zinc, this is incidental. In principle,the advantage of the invention would be obtained were the smeltingcharge made up of a mixture of lead-free franklinite and zinc-freeroasted galena.

Where the zinc smelter does not practice sulphidesintering, but roaststhe sulphide ore to form oxidized zinc compounds, the roasted ore may becoke-sintered and the resulting sinter smelted in accordance with theprinciples of the invention. A suitable coke-sintered product forsmelting in accordance with the invention may be made by sintering acharge of about the following composition:

Parts Roasted ore 57 Recirculated material (dross etc.) 12 Limestone 23Sand 6 Coke 2 The aim in coke-sintering is to obtain the whole of thesinter cake in a form suitable for smelting in a blast furnace withoutrecycling the sintered product. Some fine material, however, is alwaysobtained, usually in small amount, and this is returned for sinteringwith another charge of roasted ore.

It will be clear to those skilled in this art that the above data is byway of illustrating the invention, and that the invention lends itselfto various modifications in practice.

We claim:

1. In the method of operating a zinc blast furnace in which a preheatedcharge containing oxidic zinc materials and carbonaceous fuel isintroduced into the top of the furnace, preheated air is blown into thebottom of the furnace, molten slag is tapped from the bottom of thefurnace, a gas containing zinc vapour, carbon monoxide and carbondioxide is withdrawn from a higher part of the furnace and passed into acondenser to recover molten zinc, the improvement which comprisesincorporating a substantial amount of oxidic lead compounds in thecharge and regulating the amount of combustible carbon charged so thatthe heat generated by combustion of this carbon in the preheated air toproduce a mixture of carbon monoxide and carbon dioxide in such a ratiothat this gaseous mixture is highly reducing to the oxidic zinc and leadcompounds present in the slag formed in the furnace, together with theheat generated by the exothermic reaction between the oxidic leadcompounds and carbon monoxide to produce lead metal and carbon dioxide,is substantially no more than that necessary (a) for compensating forheat losses from the furnace, (b) for reducing the oxidic zinc compoundsendothermically by carbon monoxide to produce zinc vapour and carbondioxide, for melting substantially all of the charge components otherthan zinc oxide, lead oxide and carbon, and maintaining the resultantslag at such an elevated temperature that substantially all the zinc andlead oxides dissolved therein are reduced to liberate zinc vapour andlead metal, (at) for raising the temperature of the gases in the zone ofthe furnace where they last make contact with the furnace charge to apoint at which the zinc vapour contained therein cannot react with thecarbon dioxide contained therein to form undesired zinc oxide, and (e)for volatilizing some of the lead metal formed, the amount of leadvolatilized not being substantially above that unavoidable minimumamount required to saturate the gases leaving the charge with leadvapour at the temperature at which these gases finally leave the charge;the gases being brought into intimate shock-chilling contact with moltenlead in the condenser to facilitate condensation of the zinc vapour tomolten zinc and withdrawing molten lead from the bottom of the furnace.

2. Method according to claim 1, in which the sulphur content of thecharge does not exceed about 1.5%.

3. Method according to claim 1, in which the sulphur content of thecharge is less than 0.8%.

4-. Method according to claim 1, in which the lead content of the chargeis substantially in excess of 5% of the weight of carbon burned.

5. Method according to claim 1, in which free lime is present in thecharge to react with a substantial portion of the sulphide compoundspresent in the charge and gases to form calcium sulphide and therebysubstantially inhibit the volatilization of lead sulphide.

6; Method according to claim 1, in which the sulphur content of thecharge does not exceed about 1.5%, and

free lime is present in the charge to react with a substantial portionof the sulphide compounds present in the charge and gases to formcalcium sulphide and thereby substantially inhibit the volatilization oflead sulphide.

7. Method according to claim 1, in which the sulphur content of thecharge is less than 0.8%, and free lime is present in the charge toreact with a substantial portion of the sulphide compounds present inthe charge and gases to form calcium sulphide and thereby substantiallyinhibit the volatilization of lead sulphide.

8. Method according to claim 1, in which the temperature of the gaseswhere they last make contact with the charge is only slightly above theequilibrium temperature below which the zinc vapour could react with thecarbon dioxide to form Zinc oxide thereby inhibiting the passage of leadsulphide to the condenser.

9. Method according to claim 1, in which the gases are withdrawn fromthe top free level of the charge, where they are in contact with freshlyintroduced charge material, and the temperature of the gases where theylast make contact with the top of the charge is only slightly above theequilibrium temperature below which the zinc vapour could react with thecarbon dioxide to form zinc oxide thereby inhibiting the passage of leadsulphide to the condenser.

10. Method according to claim 1, in which a controlled amount of anoxygen-containing gas is introduced directly into the gaseous productwithdrawn from the charge, and the gaseous product is heated by theresulting oxidation of carbon monoxide to a temperature to ensure thatno appreciable amount of the zinc vapour reacts with carbon dioxideduring the passage of the gaseous product to the condenser.

11. Method according to claim 1, in which the temperature of the gaseswhere they last make contact with the charge is only slightly above theequilibrium temperature below which the zinc vapour could react with thecarbon dioxide to form zinc oxide, the passage of lead sulphide to thecondenser thus being inhibited; a controlled amount of oxygen containinggas is introduced directly into the gases so Withdrawn; and the gasesare heated by the resulting oxidation of carbon monoxide to atemperature to ensure that no appreciable amount of the zinc vapourreacts with carbon dioxide during the passage of the gases to thecondenser.

12. Method according to claim 1, in which the zinc content of themetalliferous material is derived for the most part from the inclusiontherein of a lead blast-furnace slag containing at least 10% of zinc.

13. Method according to claim 1, in which the charge includes as itsmain metalliferous components zinc-containing lead blast-furnace slagand sinter-roasted lead sulphide ore.

14. Method according to claim 1, in which dross from the condenser issintered, and the resulting sinter is added to the charge.

15. Method according to claim 1, in which blue powder obtained byscrubbing the gases coming from the condenser is sintered, and theresulting sinter is added to the charge.

16. Method according to claim 1, in which copper is present in thecharge, molten lead accumulates in the bottom of the furnace in amountto dissolve the copper, and the copper is tapped in solution with thelead.

17. Method according to claim 1, in which silver is present in thecharge, molten lead accumulates in the bottom of the furnace in amountto dissolve the silver, and the silver is tapped in solution with thelead.

18. In the method of recovering zinc by a blast furnace operation inwhich a preheated charge containing oxidic zinc materials andcarbonaceous fuel is introduced into the top of the furnace, preheatedair is blown into the bottom of the furnace, molten slag is tapped from'the bottom of the furnace, a gas' containing zinc vapor, carbonmonoxide and carbon dioxide is withdrawn from References Cited in thefile of this patent UNITED STATES PATENTS 487,444 Hunicke Dec. 6, 18921,072,209 Desgraz Sept. 2, 1913 1,652,184 Skogmark Dec. 13, 19271,949,905 Hall Mar. 6, 1934 FOREIGN PATENTS 5,438 Great Britain 1913

18. IN THE METHOD OF RECOVERING ZINC BY A BLAST FOR NACE OPERATION INWHICH A PREHEATED CHARGE CONTAINING OXIDIC ZINC MATERIALS ANDCARBONACEOUS FUEL IS INTRODUCED INTO THE TOP OF THE FURNACE. PREHEATEDAIR IS BLOWN INTO THE BOTTOM OF THE FURNACE, MOLTEN SLAG IS TAPPED FROMTHE BOTTOM OF THE FURNACE, A GAS CONTAINING ZINC VAPOR CARBON MONOXIDEAND CARBON DIOXIDE IS WITHDRAWN FROM A HIGHER PART OF THE CHARGE ANDPASSED INTO A CONDENSER TO RECOVER MOLTEN ZINC, THE IMPROVEMENT WHICHCOMPRISES SEPCIALLY ADDING A SUBSTANTIAL AMOUNT OF OXIDE LEAD COMPOUNDSTO THE CHARGE, AND USING SUBSTANTIALLY NO MORE CARBONACEOUS FUEL IN THECHARGE THAN THAT NORMALLY REQUIRED TO REDUCE THE ZINC OXIDE AND TOCONDENSE AND RECOVER THE MOLTEN ZINC SAID CARBONACEOUS FUEL BEINGSNIFFICIENT ALSO TO REDUCE THE LEAD OXIDE TO LEAD METAL, AND SEPARATELYRECOVERING MOLTEN LEAD FROM THE LOWER PART OF THE CHARGE.